Gold and silver recovery from polymetallic sulfides by treatment with halogens

ABSTRACT

A method for treating a polymetallic sulfide ore containing gold and/or silver, and further containing base metals selected from the group consisting of iron, aluminum, chromium, titanium, copper, zinc, lead, nickel, cobalt, mercury, tin, and mixtures thereof, is disclosed. The method comprises the steps of grinding the polymetallic sulfide ore to produce granules, oxidizing the granules to produce oxidized granules, and chloride leaching the granules using a brine solution including dissolved halogens, as well as chloride and bromide salts.

[0001] This application claims the benefit of U.S. ProvisionalApplication No. 60/446,517, filed Feb. 12, 2003. The entire text of theabove provisional application is specifically incorporated by reference.

FIELD OF THE INVENTION

[0002] The present invention relates to gold and silver recovery frompolymetallic sulfides by treatment with halogens.

BACKGROUND OF THE INVENTION

[0003] The use of chemical agents, particularly halides, for therecovery of gold and silver is well known. It was noted very early thatthe adjunction of sodium chloride to mercury improved the performancesof the amalgamation process. This discovery translated into the Patio orCazo processes, which were implemented on an empirical basis from theearly 1600's in Central and South America more than 150 years before thediscovery of elemental chlorine by Scheele in 1774. The Patio methodinvolved the digestion of a finely divided gold ore with mercury andsodium chloride, in the presence of air and moisture over a three monthperiod. The values were then collected by further leaching with mercury,followed by amalgam distillation (Egleston, 1887).

[0004] Later, in the late 1700s, chloridizing roasting followed bybarrel amalgamation was developed in Central Europe as an improvedmethod for gaining access to precious metals from sulfide ores. Thisprocess called upon a high temperature treatment of the gold/silver oresin the presence of sodium chloride, air and steam, in order to transformthe precious metal sulfides into their corresponding chlorides. The goldand silver was then recovered either by amalgamation or cementation onpure copper (Varley et al, 1923). However, it was discovered that thehigh temperature chloridizing of gold or silver ores resulted in veryimportant losses of values by volatilization. In some cases these lossesreached 80% or more of the precious metal content (Christy, 1888).

[0005] It appeared that the presence of pyrites or iron sulfidescontributed significantly to the volatilization of gold and silverduring high temperature chloridization with NaCl (Croasdale 1903). Itwas finally established that the mechanism explaining these lossesinvolves the formation of a mixed chloride of gold and iron(AuCl₃.FeCl₃), which is highly volatile at chloridization temperatures(Eisele et al.).

[0006] Elemental chlorine dissolved in water, introduced by Plattneraround 1850, constituted an alternative to high temperaturechloridization. However, this process was characterized by lowefficiency.

[0007] The general characteristics of the various processes involvingchlorine, either as elemental chlorine or as chlorides, either atambient temperatures or at high temperatures, were not attractive. Theyields obtained with these processes were generally low (often below50%) and the values were collected as amalgams or as cemented productson copper or iron. In addition, complex procedures were involved inorder to obtain the precious metals in a pure form. The environmentalimpacts of such operations, where large amounts of sulfur are disposedwith the tailings, would have been completely unacceptable by currentstandards.

[0008] The advent of cyanide extraction in 1916, terminated theextraction of gold by various forms of chloridation. The cyanide processcalls upon the action of a cyanide salt such as sodium cyanide on goldin the presence of oxygen, to give a soluble gold salt (Eq. I):

2Au+4NaCN+1/2O₂+H₂O→2Na[Au(CN)₂]+2NaOH   (Eq. I)

[0009] The gold can then be recovered from the cyanide complex by theaction of excess zinc (Eq. II):

2Na[Au(CN)₂]+Zn_((excess))→Na₂[Zn(CN)₄]+2Au   (Eq. II)

[0010] Under the best circumstances, gold recovery can be as high as98%. This process calls for a contact time of one to three days at nearambient temperature in the presence of air.

[0011] In some instances the cyanide process performs very poorly. Oresrefractory to cyanide extraction can be grouped under the general termof polymetallic ores. In such ores, one finds small amounts of basemetals such as copper or zinc, typically 0.1% Cu or 0.3% Zn. Such smallamounts qualify the ore as of very low grade for the production ofcopper or zinc. If such a polymetallic ore body contains some gold (forexample, 4 g/T Au or Ag or a mixture of both), the cyanide extractionprocess does not perform well. The poor performance is due to the basemetals, either copper or zinc, (as well as silver), having a muchstronger ability to form complexes with cyanide than gold. In fact, thisinherent property is used to recover gold from a pregnant solution byzinc treatment following cyanide extraction (see Eq. II). The basemetals will consume all the cyanide present and the gold extraction willonly begin after all the available base metals, as well as silver, havebeen dissolved. Because of the excessive consumption of relativelycostly cyanide, this process for recovering gold is uneconomical.

[0012] Polymetallic ores constitute complex mixtures of sulfides. Thetailings discarded as a result of gold and silver extraction using thecyanide process, as well as by other methods, still contain verysubstantial amounts of sulfur. This sulfur is prone to bio-oxidation(Thiobacillus ferrooxidans), and the resulting drainage is quite acidicand toxic due to its metallic content.

[0013] The spent cyanide solutions, kept in large ponds following goldrecovery, represents a substantial environmental hazard and has recentlycreated major disasters in Guyana and Central Europe, thus restrictingthe use of the cyanide process in many areas.

[0014] In the last twenty years, chloridation has been reconsidered as aprocess for extracting base metals such as copper, nickel or silver. TheIntec Base Metal Process (Moyes and Houllis, 2002) constitutes a typicalexample. This process calls for the digestion at 85° C., over a periodranging from 12 to 14 hours, of the sulfides of copper or zinc in aconcentrated brine solution (250 g/l NaCl) comprising a cupric mixedhalide (BrCl₂)Cu prepared electrolytically. The mixture is aerated andthe copper is collected as cuprous chloride. The cuprous chloride isdecomposed at the cathode to elemental copper by electrolysis uponregeneration of the mixed halide of copper (Eq. III):

2CuFeS₂+5BrCl₂ ⁻→2Cu⁺²+2Fe⁺³4S^(o)+5Br⁻+10Cl⁻  (Eq. III)

[0015] The above described chloridation process was reported as alsoextracting gold, if present. However, the requirement of recyclingcopper so as to have the cupric/cuprous system needed to oxidize ironand sulfur, makes this approach very cumbersome when the main concern isgold recovery rather than copper recovery. Further, the electrolyticaloxidation of sulfur via the cupric salt, which is regenerated byelectrolysis, is a very costly process rendering the treatment of a goldore having a modest gold content uneconomical. Finally, the presence ofelemental sulfur in the tailings is a potential source of acid drainage.

[0016] Another chloridation process called Platsol, was reported asbeing very efficient for the recovery of base and precious metals fromsulfide ores (Ferron et al, 2002). This process involves a pressureoxidation in the presence of oxygen and sulfuric acid in an autoclave ata temperature above 200° C. The implementation of such a technique isvery capital-incentive, calling for titanium autoclaves and a source ofpure oxygen. The operation of this equipment is also prone to problemsdue to scaling of the reactor, complicating heat transfer. The sulfurresulting from the operation is in an innocuous form, i.e. a hydratediron sulfate (jarosite). The high capital and operating costs renderthis approach unattractive for polymetallic sulfides having a modestgold content.

[0017] Other techniques such as the Plint process (Frias et al, 2002)or, the Ito process (Kappes et al, 2002), are techniques used for therecovery of gold and silver from sulfides, by oxidation with ferricchloride in concentrated brine. The ferrous chloride is re-oxidized toferric chloride by chlorine alone or by exposure to air and hydrochloricacid (Eq. IV):

2PbS.Ag₂S.3Sb₂S₃+24FeCl₃→24FeCl₂+2PbCl₂+2AgCl+6SbCl₃+12^(o)   (Eq. IV)

[0018] In these processes, sulfur is again oxidized electrochemicallyvia the oxidation of ferrous chloride by chlorine or HCl. As explainedpreviously, such an approach is costly for the recovery of gold orsilver from sulfide ores, because of the electrochemistry involved.Elemental sulfur is again discarded with the tailings, generating apotential source of acid drainage.

[0019] There thus remains a need for an improved method for the recoveryof gold and silver from polymetallic ores.

[0020] The present invention seeks to meet these and other needs.

SUMMARY OF THE INVENTION

[0021] The present invention relates to a method for treating apolymetallic sulfide ore containing gold and/or silver, and furthercontaining base metals selected from the group consisting of iron,aluminum, chromium, titanium, copper, zinc, lead, nickel, cobalt,mercury, tin, and mixtures thereof, comprising the steps of:

[0022] (a) grinding the polymetallic ore to produce granules;

[0023] (b) oxidizing the granules at temperatures of at least about 300°C. to produce oxidized granules;

[0024] (c) chloride leaching the oxidized granules to produce a pregnantsolution of solubilized metal chlorides and a barren solid;

[0025] (d) recovering the barren solid from the pregnant solution toproduce a purified pregnant solution; and

[0026] (e) selectively recovering gold and/or silver from the purifiedpregnant solution yielding a solution essentially deprived of goldand/or silver.

[0027] The present invention further relates to a method for therecovery of gold and silver from polymetallic sulfide ores,characterized by low operational and cost investments.

[0028] The present invention also relates to a method for the recoveryof gold and silver from polymetallic sulfide ores, characterized bybeing carried out at atmospheric pressure and at low oxidationtemperatures prior to leaching.

[0029] In addition, the present invention relates to a method for therecovery of gold and silver from polymetallic sulfide ores,characterized by producing tailings devoid of elemental sulfur,sulfides, or soluble sulfates and by fast reaction rates allowing forhigh rates of treatment.

[0030] Furthermore, the present invention relates to a method for therecovery of precious metals such as gold and silver, as well as basemetals such as copper, nickel, cobalt, zinc, tin and lead frompolymetallic sulfide ores, in addition to relating to the safe removalof sulfur, arsenic and mercury as well as to the disposal of iron,chromium, aluminum and titanium in an inert and insoluble form.

[0031] Further scope and applicability will become apparent from thedetailed description given hereinafter. It should be understood however,that this detailed description, while indicating preferred embodimentsof the invention, is given by way of illustration only, since variouschanges and modifications within the spirit and scope will becomeapparent to those skilled in the art.

BRIEF DESCRIPTION OF THE DRAWINGS

[0032] In the appended drawings:

[0033]FIG. 1 is a block diagram illustrating the various steps of themethod of the present invention;

[0034]FIG. 2 is a block diagram illustrating the various steps of thesulfur removal aspect of the method of the present invention;

[0035]FIG. 3 is a block diagram illustrating the various steps of thegold and silver recovery aspect of the method of the present invention;and

[0036]FIG. 4 is a block diagram illustrating the various steps of thebase metal recovery aspect of the method of the present invention; and

[0037]FIG. 5 is a schematic illustration of an electrolytic cell used inthe method of the present invention.

DETAILED DESCRIPTION OF THE INVENTION

[0038] Unless defined otherwise, the scientific and technological termsand nomenclature used herein have the same meaning as commonlyunderstood by a person of ordinary skill. As defined herein, theterminology “recovering” is understood as being an operation resultingin the separation of a solid from a liquid. Non-limiting examples ofsuch an operation include filtration techniques such as gravityfiltration, pressure filtration, vacuum or suction filtration andcentrifugation.

[0039] In a broad sense, the present invention relates to a new methodfor the recovery of precious metals such as gold and silver frompolymetallic sulfide ores. In an other aspect, the present inventionalso relates to the safe removal of sulfur, arsenic and mercury as wellas to the disposal of iron, chromium, aluminum and titanium in an inertand insoluble form. This is achieved at considerably lower cost thanwith the current chloridation or cyanide processes, by avoiding sulfuroxidation by electrochemical means. The method of the present inventionis very time efficient, of the order of a few hours, and is carried outat atmospheric pressure and at oxidation temperatures of at least about300° C. and preferably ranging from about 400 to about 600° C. Themethod allows for the separation of the precious metals as well as thebase metals from the common metals, while recycling the reagents andreleasing only inert waste materials into the environment.

[0040] In a preferred embodiment, gold and silver, and optionally basemetals such as copper, zinc, lead, tin, nickel, cobalt and mercury canbe recovered from polymetallic sulfide ores in yields generally wellabove 80% by the method of the present invention comprising thefollowing preferred steps:

[0041] oxidizing the polymetallic sulfide ore, preferably using lean airhaving about 10% O₂, at a temperature ranging from about 400 to about600° C., to reduce the sulfur content of the ore to about 0.5% S (assulfide) or less. Temperatures above 600° C. are also suitable butenergy consumption is increased and sintering of the ore results. Theresulting SO₂ is fixed by calcium carbonate as calcium sulfite, whichauto-oxidizes to calcium sulfate dihydrate (gypsum). This results in theelimination of sulfur in a manner compatible with environmentalregulations;

[0042] leaching the sulfur-free ore with a near-saturated (275 to 300g/l) aqueous solution of sodium chloride (sodium brine), or a nearsaturated (190 to 225 g/l) aqueous solution of potassium chloride(potassium brine) and adding a solution of chlorine/HCl/hypochlorousacid such that the precious metals and the base metals are chlorinatedand dissolved in the strongly complexing brine milieu. The chloridationreaction is advantageously and significantly accelerated by thepreferred presence of a catalytic amount, less than one percent of thehalides present in the brine, of bromide ions. Thechlorine/HCl/hypochlorous acid solution, containing a catalytic amountof bromine, is generated by circulating a portion of the brine solutionused to slurry the oxidized ore through the anodic compartment of anelectrolytic cell, at a rate sufficient to dissolve the chlorine in thebrine solution. Following the slurring operation, the ore is maintainedin suspension in the acidic halogenated brine at a temperature rangingfrom about 35-45° C. by slow stirring, without aeration, for a period of2-3 hours for most ores, and up to 5 hours for exceptionally refractoryores. After separating the barren solid followed by washing with brine,the combined filtrate and rinsings are circulated over activated carbonfor gold and silver recovery; and

[0043] treating the solution deprived of precious metals with a sodiumhydroxide solution (or a potassium hydroxide solution if potassium brinewas used) raising the pH to about 2.5-3.5. The sodium hydroxide (orpotassium hydroxide) required to achieve this partial neutralization isproduced by circulating the initial brine solution through the cathodiccompartment of the electrolytic cell. The caustic sodium hydroxidesolution (or potassium hydroxide solution) is generated concomitantly atthe cathode, in a stoïchiometric ratio, with the chlorine/hydrochloricacid/hypochlorous acid solution produced at the anode of theelectrolytic cell. Raising the pH to about 2.5-3.5 induces theprecipitation of iron, aluminum, chromium and titanium as insolubleoxides of these metals, in various hydrated forms. These oxides arefiltered and washed with brine. Raising the pH of the resulting filtrateto values above 3.5, induces the precipitation of the base metals suchas copper, zinc, lead, tin, nickel and cobalt as a base metalconcentrate.

[0044] Any arsenic, often present in significant amounts in polymetallicsulfide ores, is eliminated along with the sterile solids followingleaching as ferric arsenate, an insoluble and inert arsenic salt.Mercury, if present, is largely recovered with the flue dusts afteroxidation, and any remaining traces of this metal are lixiviated by thechlorinated brine, and recovered on carbon together with gold andsilver.

[0045] The brine solution, following the removal of the metals, isrecirculated for further leaching. The sterile solids are rinsed withwater and the rinsings concentrated by evaporation, using waste heatfrom the sulfide oxidation step. The concentrated rinsings, along withthe brine solution, are then recycled so as to prevent salt losses orsalt release into the environment.

[0046] Sulfur Removal (FIG. 2)

[0047] The gold and/or silver containing ore, additionally comprisingvariable amounts of base metals such as Cu, Zn, Pb, Sn, Ni, and Co, is asulfide or complex sulfide. The ore may further incorporate one or moreother common metals such as iron, aluminum, titanium, chromium, as wellas elements such as arsenic, antimony or bismuth. Mercury isoccasionally also present in the ore.

[0048] The ore is reduced to a particle size of less than about 140 meshby standard methods known in the art, such as crushing. The sulfurcontent of the ore, which can be as high as 15%, is reduced to about0.5% or less (as sulfides) by controlled oxidation in a reactor or kiln.The reactor or kiln provides for a control of the oxygen content in thereaction chamber. A relatively low oxidation temperature, typicallyranging from about 400 to about 600° C., is very advantageous since itprevents any sintering of the material and generates a solid producthaving a large surface area and having good reactivity. This treatmentis much preferred to standard roasting where temperatures as high as1200° C. have been observed. Such high reaction temperatures induce muchsintering and volatilization. Standard roasting involves the freeburning of the sulfides in the presence of excess air.

[0049] The control of the low oxidation temperatures is achieved byrecycling part of the lean air back to the reactor. This allows for theoxygen content in the reactor to be maintained at values not exceeding10% O₂. It is important to prevent sodium chloride present in the orefrom being oxidized. It is well known that sodium chloridecontaminations as low as 0.01 percent, can induce significantvolatilization of gold and silver.

[0050] The gas stream from the oxidation reactor is cooled in a settlingchamber, allowing for the collection of volatile oxides such as arsenicoxide, traces of zinc oxide, and metallic mercury if present in thestarting ore, as well as other products generated during the oxidativetreatment. Dusts carried mechanically from the fines in the reactor arealso collected in the settling chamber. The amount of solids collectedis generally small; less than one percent of the weight of the oretreated. The solids thus collected can be recovered and used forrecuperation of values such as As₂O₃ or mercury, or they can be safelydisposed of in sealed containers. The gas at the exit of the settlingchamber, essentially composed of SO₂ and lean air, is partly redirectedback to the oxidation reactor for oxygen level control, and partlydirected to a SO₂ scrubbing unit. The SO₂ is adsorbed using a finelydivided limestone slurry (200 mesh), allowing for the transformation ofessentially all of the SO₂ (about 98%) into calcium sulfite, whichauto-oxidizes to calcium sulfate dihydrate or gypsum. Gypsum is a verystable and inert product representing a definitive solution for the safedisposal of sulfur. It can be used as a building material in theproduction of Portland cement or as land fill. The water following thedewatering of the gypsum is recirculated back to a water thank. Sincegypsum is a dihydrate, there is a net consumption of water in thescrubbing process. The gases freed of SO₂, are vented through a flueduct.

[0051] In the first step of the method therefore, the ore was made morereactive towards leaching, and essentially all of the sulfur initiallypresent has been disposed of in a safe and environmentally compatiblemanner. The present approach constitutes an economically attractivealternative to the presently available methods. The current cost ofelectrochemically oxidizing 1% of sulfur in one metric ton of sulfideore is $US 4.71 per unit percent of S²⁻ per ton with a KWh at $US 0.09per kilowatt and with an efficiency of 80%. The cost of oxidizing thesulfide content of an ore containing 10% S²⁻ to elemental sulfur, usingan electrochemically-produced reagent such as chlorine, would be in thebest case scenario $US 47.10 per ton of ore for power only. Thecontrolled oxidation of the sulfur content using lean air can be done at10% or less of that cost, and transforms the sulfur into a safe andenvironmentally disposable form. The electrochemical oxidation processleaves elemental sulfur in the tailings generating a potential source ofacid drainage.

[0052] Gold/Silver Recovery (FIG. 3)

[0053] The recovery of gold and silver from the oxidized ore is achievedby leaching with a reagent comprising elemental halogens. The halogens(Br₂, Cl₂) have significantly different behaviors towards gold. Brominecan readily dissolve gold at room temperature, even in the absence ofwater (Kruss and Schmidt, 1887). Gold, on the other hand, is inert todry chlorine at room temperature, and the attack of this gas on goldrequires the presence of water and slight heating (Voigt and Biltz,1924). Even though bromine is the more reactive reagent with gold,chlorine is more electronegative (Latimer, 1952):

Cl⁻→Cl₂(−1.359V);

Br⁻→Br₂(−1.07V).

[0054] It is possible to take advantage of this reactivity difference toaccelerate gold leaching from the oxidized ore, if a catalytic amount ofa bromide is introduced into the leaching solution. The leachingsolution is a brine solution having a high concentration of chloride,i.e. from 275 to 300 g/l of NaCl or from 190 to 225 g/l of KCl. Lowersalt concentrations yielded lower percentages of silver recovery, whensilver was associated with gold in the oxidized ore. A portion of theconcentrated brine solution, also containing a trace (1-3 g/l) of NaBror KBr, is circulated in the anodic compartment of an electrolytic cell,at an appropriate rate, so as to dissolve the halogen liberated at theanode. As mentioned above, the bromide ion will be reduced first,followed by some chloride ions so as to give a mixture of halogensdissolved in the brine solution. The brine solution containing dissolvedCl₂ and Br₂ is mixed with fresh brine from a brine tank to provide avolume of liquid necessary to form a 20% slurry with the oxidized ore ina reactor kept at 35-45° C. The slurry is slowly stirred in order toprevent settling of the ore. The reacting mass was not aerated sinceaeration was neither improving the reaction rate nor the reaction yield,instead it resulted in the loss of dissolved halogens. Due to the traceamounts of bromine in the system, the gold leaching process is believedto involve the initial formation of gold tribromide (Eq. V):

2Au+3Br₂→2AuBr₃   (Eq. V)

[0055] The gold tribromide is then believed to be transformed, becauseof the stronger oxidizing capacity of Cl₂, into gold trichloride withthe concomitant regeneration of elemental bromine (Eq. VI):

2AuBr₃+3Cl₂→2AuCl₃+3Br₂   (Eq. VI)

[0056] A similar type of reaction is obtained for silver, the highconcentration of chloride allowing the solubilization of the silverhalides by complexation.

[0057] In the course of the leaching reaction, the other ions aresimilarly solubilized, and exist at their maximum valency; copper ascupric chloride, iron as ferric chloride, tin as stannic chloride, andarsenic as arsenate (As⁺⁵). Particularly with arsenic, the strongoxidizing environment leads to the precipitation of all the arsenic asan insoluble and inert ferric arsenate (Eq. VII):

Fe³⁺+AsO₄ ⁻³→FeAsO₄   (Eq. VII)

[0058] The pH of the reaction mixture drops below 0.1 as the leachingreaction proceeds. This strong acidification is an indication of thereaction of chlorine with water (Eq. VII):

H₂O+Cl₂→HCl+HOCl   (Eq. VIII)

[0059] The presence of hypochlorous acid could account for the observedchloridation of gold by chlorine in the presence of water. A similarequation can be written to describe the behavior of bromine, which is inequilibrium with hydrobromic acid and hypobromous acid. The solubilizedspecies can therefore be seen as a mixture of chlorides andhypochlorides, which eventually end up as chlorides when thehypochlorous ion decomposes with the concomitant evolution of nascentoxygen (Eq. IX):

HOCl→HCl+1/2O₂   (Eq. IX)

[0060] The production of nascent oxygen accounts in part for the verystrong oxidizing capability of the system without aeration of any sort.

[0061] The duration of the leaching, preferably at 35-45° C. in thereactor, usually ranges from 2 to 3 hours. With exceedingly refractoryores it is necessary to extend the contact time to, for example, about 5hours. Following the leaching, the slurry is filtered or centrifuged,producing a pregnant solution and a waste or barren solid.

[0062] The barren solid was first rinsed with brine in order to recoverany held-up values in the cake, followed by washing with water torecover any salt. The so-obtained tailings contain arsenic as an ironarsenate, and are free of sulfur and of soluble base metals. Thepregnant solution is circulated over carbon to collect the gold andsilver. Following the recovery of gold and silver from the carbon byknown methods, these precious metals are obtained by electrowinning orother standard techniques such as ion exchange and precipitation. Thegold/silver-free solution is then recovered to be further treated so asto collect the base metals.

[0063] Recovery of Base Metals (FIG. 4)

[0064] The base metals to be obtained from the leaching of gold-bearingpolymetallic sulfide ores are of two categories. The first categorycontains metals of relatively high commercial value, often obtained bypyrometallurgical operations. This category contains metals such asnickel, cobalt, copper, zinc, lead, tin and mercury. The second categorycontains metals of low economic value, and comprises predominantly ironwith considerably smaller amounts of aluminum, titanium, chromium andtraces of the p-bloc elements.

[0065] In order to isolate these two types of base metals, sodiumhydroxide is generated in the cathodic compartment of the electrolyticcell. The sodium hydroxide solution is accumulated in a caustic tank andis then used to raise the pH of the previously produced barren solution,devoid of gold and silver, leaving the carbon columns, from below 1 toabout 2.5 to about 3.5. At a pH ranging from about 2.5 to about 3.5, anyiron existing as Fe⁺³ is instantaneously precipitated by hydrolysis as ahydrated iron oxide. Titanium, aluminum and chromium react similarlywithin this pH range. The hydrated oxides are removed by filtration. Thesolids are rinsed with brine in order to recuperate any base metals ofvalues held up in the solid cake, followed by washing with water toremove any traces of salt. The salt-free mixture of oxides is thendiscarded as an insoluble and inert material of little or no commercialvalue.

[0066] The solution obtained from the filtration and the brine rinsingscontains the base metals of value. Mercury, if present, was recovered oncarbon together with gold and silver. The pH of the mercury-freesolution, pH between about 2.5-3.5, is further raised using anadditional portion of the sodium hydroxide solution to values above 3.5,causing all of the base metals (Ni, Co, Cu, Zn, Pb, Sn) to precipitateas oxides or hydrated oxides. The oxides are removed from the mixture byfiltration and are rinsed with water to remove any traces of salt, toprovide a concentrate of metals having significant commercial value. Thebrine, being free of metals, is recycled back to the fresh brinereservoir. The rinsings are concentrated by evaporation so as to give abrine solution of appropriate concentration, and which is also recycledback to the fresh brine reservoir.

[0067] The implementation of the process of the present invention, usinga large variety of gold-bearing polymetallic sulfide ores, provides forthe recovery of gold and silver in high yields, essentially always above80% and frequently above 85%. The process of the present invention alsoprovides for the recovery in high yields of the base metals ofcommercial value, frequently above 85%.

[0068] Of all the base metals of little commercial value, iron isgenerally the predominant one. Following the oxidation of the sulfidesat 400-600° C., the resulting iron oxide is quite inert and no more thanabout 20-25% of this iron is leached, thus significantly decreasing thepower consumption of the process. In fact, for a KWh costing US$ 0.09,and with an efficiency at the electrolytic cell of 80%, each percent ofiron in the ore would cost US$ 1.00 of power to take care of, and eachpercent of base metals such as copper or zinc in the ore would cost US$2.36 of power to extract. Thus, for an ore having 1% copper and 8% iron,the value of recovered copper (US$ 16.50 at US$ 0.75/lb for copper)covers all the electrolytical power costs (US$ 10.36) plus a fairreserve and no power imputations have to be made against the gold andsilver values recovered.

[0069] Using the process of the present invention, polymetallic sulfideores containing gold and/or silver which do not qualify for base metalsextraction either because of a low base metal content, problems ofenrichments by flotation or other restrictions, can be treatedeconomically from the return generated by the base metals in order tocollect the precious metals. Consequently, the process of the presentinvention provides for an attractive alternative to the currentlyavailable technologies, allowing the treatment of ores or tailingspreviously not attractive, at a profit.

[0070] The recycling of the brine solution, and the disposal of sulfur,arsenic and metal oxides as stable and inert solids, reduces theenvironmental impacts of the operation to a minimum. Furthermore, theimplementation of the process of the present invention at low oxidationtemperatures, at near ambient chloridation temperatures and atatmospheric pressure, reduces the investment per unit weight of ore tovery competitive values. Finally, the low temperature oxidation ofsulfur being an exothermic process, the energy consumption at that levelis minimal and much lower than the corresponding electrochemicaloxidation of sulfide to elemental sulfur.

[0071] The process of the present invention was tested using a varietyof polymetallic sulfide ores and tailings containing gold and silver.

EXAMPLES Example 1

[0072] A Canadian ore sample (90 g) from the Sudbury (Ontario) areacontaining 4.5 g/T Au, 8 g/T Ag, 0.1% As, 7.5% S, 5.5% Fe, 0.1% Ni,0.008 Co and 0.5% Cu was reduced to a particle size of about 140 meshand heated at 585-600° C. in an atmosphere composed of N₂ (50%) and air(50%), over a period of two hours in a Vycor™ tube heated externally ina Lindberg™ furnace. The temperature was measured inside the mass beingoxidized. The external heating was reduced when the oxidation began ataround 400° C.

[0073] A small amount of a white deposit, arsenic oxide, could beobserved at the discharge side of the Vycor™ tube. The color of theoxidized material changed from black to brown and the weight loss duringthe process was about 12%.

[0074] A sample of the oxidized material (25.0 g) was placed in athree-necked one liter flask, along with 500 g of water, 150 g of sodiumchloride and 1.2 g of sodium bromide. The suspension was stirredmagnetically and the flask was closed so as to exclude air from enteringthe system.

[0075] The slurry was extracted from the flask through one of the necksusing a peristaltic pump, and was subsequently circulated through theanodic compartment of an electrolytic cell operating with a brinesolution having the same concentration as the brine solution in theflask (anode of graphite, operation at 2.5 V). The anodic fluid wasreturned to the flask after dissolving chlorine. The cell was operatedon and off in such a manner as to maintain a slight reddish colorationin the flask indicative of the presence of free bromine.

[0076] The reaction flask was maintained at 40° C. for a period of 2.5hours after which it was filtered on a Buchner funnel. The solid wasrinsed three times with a brine solution containing 300 g/l NaCl. Themixed filtrate and rinsings were very acid, having a pH below 1.0. Theacidic filtrate and rinsings were then treated with 30 g of carbon(Norit™ RO3515) so as to collect gold and silver. The barren solid wasthen rinsed with water to completely remove any traces of brine(negative test to AgNO₃), dried at 110° C. (16.8 g) and submitted toelemental analysis. The elemental analysis indicated that 96% of thegold and 94% of the silver initially present in the oxidized material,were leached out and then adsorbed on the carbon.

[0077] The solution following contacting with carbon was combined withthe aqueous rinsings and was submitted to elemental analysis. Thesolution was found to be essentially free of gold and silver, andcontained 99% of the extracted iron, 98% of the nickel and copper and91% of the cobalt present in the starting oxidized ore sample. Adjustingthe pH to about 3.5 with sodium hydroxide resulted in the precipitationof the iron. Further raising the pH to about 8.5 precipitated thenickel, cobalt and copper. The brine, being essentially free of metals,is available for further use. It was noted by elemental analysis thatthe bromine content in the brine did not change during the process,taking into account the dilution induced by the rincings. Further, itwas found that the gold and silver content following treatment (in thesterile residue), was below 0.05 g/T and 0.16 g/T respectively, while23% of the iron was extracted.

[0078] The process was repeated using several types of polymetallicsulfide ores containing gold, silver or both, along with base metals ofvalue. All the operational parameters, except the duration of thedigestion, were the same as in Example 1. Those results are reported inTable I.

Example 2

[0079] A sample of ground ore (100-200 mesh) from the Pueblo Viejodeposit (100 g), Dominican Republic, and containing 3.0 g/T Au, 2.25 g/TAg, 0.28% Zn, 0.025% As, 5.8% Fe and 4.9% S (as sulfides) was oxidizedat about 600° C. for a period of 2 hours in lean air (about 10% O₂).

[0080] The oxidized material was then leached using KCl brine (50.0 g ofoxidized ore in 500 mL of KCl brine (200 g KCl/L) containing 2.0 g KBr).The suspension was stirred at 45° C. for a period of two hours, while inthe presence of chlorine (0.7 g), added to the slurry at the beginningof the contact.

[0081] The slurry was filtered, the cake rinsed with KCl brine (200 gKCl/L) and then washed with water. The combined brine filtrate, rinsingsand washings were analyzed for gold, silver and zinc. The gold recoverywas of the order of 87%; the silver recovery was of the order of 61%;and the zinc recovery was of the order of 99%. Essentially all of thearsenic was found in the barren solid, and none was present in the brineor water rinsings.

[0082] Although the present invention has been described hereinabove byway of preferred embodiments thereof, it can be modified, withoutdeparting from the spirit and nature of the subject invention as definedin the appended claims. TABLE 1 Treatment of polymetallic ores Preciousmetals content Base metals Deposit (g/T) content (%) Sulfur DurationRecovery % Ex. site Country Au Ag Cu Zn Others content % (hours) Au AgCu Zn Others 2 Zacateca*** Mexico 3.5 8.0 0.3 0.1 Pb: 0.8 7.5 2.0 94 9298 96 Pb: 91 3 Cassandra* Greece 28 12 — 1.0 Pb: 1.5 11.0 3.0 96 95 — 98Pb: 94 4 Potosi*** Bolivia 3.0 5.8 0.5 — Sn: 1.9 8.8 3.0 96 92 99 — Sn:89 5 Red Lake* Canada 17.0 16.5 0.2 0.8 — 7.3 2.5 95 96 98 — — 6Rosario*** Dom. 3.37 34.7 0.01 1.1 — 4.5 3.5 85 91 95 95 — Republic 7Moore* Dom. 5.5 8.0 0.01 1.1 — 6.0 5.0 85 88 98 99 — Republic 8 ItalianItaly 52 5100 1.13 8.06 Pb: 5.18 11.8 2.5 97 87 96 97 Pb: 99 Smelter**Hg: 1130 ppm Hg: 99.9 9 Rio Spain 231 248 25.2 0.39 Pb: 0.14 18.5 3.5 9896 99 95 Pb: 92 Narcea**

References

[0083] The following references, to the extent that they provideexemplary procedural or other details supplementary to those set forthherein, are specifically incorporated herein by reference.

[0084] Christy, Transaction of the American Institute of MiningEngineering, 17:3, 1888.

[0085] Croasdale, J. Engineering and Mining, 312, 1903.

[0086] Egleston, In: The Metallurgy of Silver, Gold and Mercury in theUnited States, 1:261, John Wiley, 1887.

[0087] Eisele et al. U.S. Bureau of Mines, Report N^(o) 7489.

[0088] Ferron et al, In: Chloride Metallurgy, Vol. I:11, CanadianInstitute of Mining, Metallurgy and Petroleum, 2002.

[0089] Frias et al, In: Chloride Metallurgy, Vol. I:29, CanadianInstitute of Mining, Metallurgy and Petroleum, 2002.

[0090] Kappes et al, In: Chloride Metallurgy, Vol. I:69, CanadianInstitute of Mining, Metallurgy and Petroleum, 2002.

[0091] Kruss and Schmidt, Berichte der Deutschen Chemichen Gesellschaft,20:2634, 1887.

[0092] Latimer, In: Oxidation State of the Elements, 56-62, PrenticeHall, 1952.

[0093] Moyes and Houllis, In: Chloride Metallurgy, Vol. II:577, CanadianInstitute of Mining, Metallurgy and Petroleum, 2002.

[0094] Varley et al, U.S. Bureau of Mines, Bulletin N^(o) 211, 1923.

[0095] Voigt and Biltz, Z. anorg. Chem., 133:277,1924.

1. A method for treating a polymetallic sulfide ore containing gold orsilver, and further comprising a base metal selected from the groupconsisting of iron, aluminum, chromium, titanium, copper, zinc, lead,nickel, cobalt, mercury, tin, and mixtures thereof, the methodcomprising: (a) grinding said polymetallic ore to produce granules; (b)oxidizing said granules at temperatures of at least about 300° C. toproduce oxidized granules; (c) chloride leaching said oxidized granulesto produce a pregnant solution of solubilized metal chlorides and abarren solid; (d) recovering said barren solid from said pregnantsolution to produce a purified pregnant solution; and (e) selectivelyrecovering gold or silver from said purified pregnant solution yieldinga solution essentially deprived of gold or silver.
 2. The method ofclaim 1, further comprising subsequent treatment of the solutiondeprived of gold or silver so as to precipitate and remove solubilizedbase metal chlorides.
 3. The method of claim 1, wherein in step (b) saidoxidizing is performed using lean air.
 4. The method of claim 1, whereinin step (c) said chloride leaching involves contacting said oxidizedgranules with a leaching solution comprising a brine solution includinga dissolved halogen.
 5. The method of claim 1, wherein in step (d) saidrecovering eliminates the barren solid from the pregnant solution ofsolubilized metal chlorides as a filtrate, and wherein the barren solidis washed with a brine solution to produce washings and a sterile solid,the washings being combined with the filtrate to produce said purifiedpregnant solution.
 6. The method of claim 2, wherein said solutiondeprived of gold or silver is treated with a caustic solution to producea first reaction mixture having a pH ranging from about 2.5 to about3.5, further producing a precipitate comprising a first set of basemetals comprising a hydrated metal oxide selected from the groupconsisting of iron, aluminum, chromium and titanium, and recovering saidprecipitate yielding a first solution essentially devoid of iron,aluminum, chromium and titanium.
 7. The method of claim 6, furthercomprising the subsequent step of treating said first solution with acaustic solution to produce a second reaction mixture having a pHranging from about 3.5 to about 14, further producing a precipitateincluding a second set of base metals comprising a hydrated metal oxideselected from the group consisting of nickel, copper, cobalt, zinc, leadand tin, and recovering said precipitate yielding a second solutionessentially devoid of nickel, copper, cobalt, zinc, lead and tin.
 8. Themethod of claim 3, wherein following said oxidizing, said lean air iscooled in a settling chamber allowing for a volatile species to becollected; wherein a first portion of said lean air and sulfur dioxideis recycled from said settling chamber to said oxidizing step; andwherein a second portion of said lean air and sulfur dioxide is directedto a sulfur dioxide scrubbing unit.
 9. The method of claim 8, wherein insaid scrubbing unit said sulfur dioxide is converted to calcium sulfatedihydrate following treatment with an aqueous limestone slurry.
 10. Themethod of claim 9, wherein said calcium sulfate dihydrate issubsequently dried.
 11. The method of claim 8, wherein said volatilespecies comprise mercury, arsenic oxide and zinc oxide.
 12. The methodof claim 8, wherein said lean air includes an oxygen content of about10%.
 13. The method of claim 3, wherein said oxidizing is performed attemperatures ranging from about 400 to about 600° C.
 14. The method ofclaim 4, wherein a first portion of a brine solution is circulatedthrough an electrolytic cell to separately and concomitantly produce acaustic solution and said brine solution including dissolved halogens,and wherein said brine solution including dissolved halogens is combinedwith a second portion of said brine solution to produce said leachingsolution.
 15. The method of claim 14, wherein said brine solutionincludes a concentration of sodium chloride ranging from about 275 g/Lto about 300 g/L.
 16. The method of claim 14, wherein said brinesolution includes a concentration of potassium chloride ranging fromabout 190 g/L to about 225 g/L.
 17. The method of claim 1, wherein instep (e) said purified pregnant solution is treated with carbon toproduce a reaction mixture including a carbon cake rich in gold orsilver, and wherein the carbon cake is removed from the reaction mixtureto produce said solution essentially deprived of gold and silver. 18.The method of claim 17, wherein said gold or silver is stripped fromsaid carbon cake and wherein said gold or silver is selectivelyrecovered by a process selected from leaching followed byelectrowinning, and precipitation.
 19. The method of claim 5, whereinsaid sterile solid is washed with water to produce a salt containingsolution, said salt containing solution being concentrated and recycledto said leaching step (c).
 20. The method of claim 6, wherein saidprecipitate is washed with a brine solution to produce washings and awashed residue, said washings being combined with said first solutionessentially devoid of iron, aluminum, chromium and titanium.
 21. Themethod of claim 7, wherein said second solution essentially devoid ofnickel, copper, cobalt, zinc, lead and tin is recycled to said leachingstep (c).
 22. The method of claim 19, wherein said salt containingsolution includes salts selected from the group consisting of sodiumchloride and sodium bromide.
 23. The method of claim 19, wherein saidsalt containing solution includes salts selected from the groupconsisting of potassium chloride and potassium bromide.
 24. The methodof claim 14, wherein said halogens are selected from the groupconsisting of chlorine and bromine.
 25. The method of claim 5, whereinsaid brine solution comprises a concentration of sodium chloride rangingfrom about 275 g/L to about 300 g/L.
 26. The method of claim 5, whereinsaid brine solution comprises a concentration of potassium chlorideranging from about 190 g/L to about 225 g/L.
 27. The method of claim 14,wherein said brine solution further includes a bromide salt selectedfrom the group consisting of sodium bromide and potassium bromide. 28.The method of claim 27, wherein said bromide salt is present in acatalytic amount.
 29. The method of claim 28, wherein said catalyticamount is ranging from about 1.0 g/L to about 3.0 g/L.
 30. The method ofclaim 4, wherein said chloride leaching is carried out at ambienttemperatures over a period ranging from about 2 to about 5 hours. 31.The method of claim 30, wherein said ambient temperatures range fromabout 35 to about 45° C.
 32. The method of claim 6, wherein said causticsolution is a sodium hydroxide solution.
 33. The method of claim 6,wherein said caustic solution is a potassium hydroxide solution.
 34. Themethod of claim 17, wherein said carbon is activated carbon.
 35. Themethod of claim 3, wherein said granules have a particle size rangingfrom about 35 mesh to about 200 mesh.
 36. The method of claim 35,wherein about 80% of said granules have a particle size of less than 35mesh and wherein about 20% of said granules have a particle size of lessthan 200 mesh.
 37. The method of claim 35, wherein about 20% of saidgranules have a particle size of less than 35 mesh and wherein about 80%of said granules have a particle size of less than 200 mesh.
 38. Themethod of claim 13, wherein said oxidized granules have a sulfur contentinferior to about 0.5%.
 39. The method of claim 19, wherein said sterilesolid may include ferric arsenate.
 40. The method of claim 1, whereinsaid gold or silver are recovered in yields in excess of about 80%. 41.The method of claim 1, wherein said polymetallic sulfide ore comprisesgold and silver.